Metal and fluorine values recovery from mineral ore treatment

ABSTRACT

The process for recovering Ta/Nb values from highly flourinated ore materials by the process of; 
     (a) contacting the materials with one or a mixture of HNO 3  or HCl, and H 3  BO 3  wherein the H 3  BO 3  to other acid molar ratio is from about 1/10 to about 1/1, 
     (b) maintaining the temperature between about 55 C. and about 85 C., 
     (c) reacting for a sufficient period to substantially solubilize the major portion of the materials and to substantially insolubilize all Ta/Nb values, and 
     (d) separating the solids containing the Ta/Nb values from solution.

FIELD OF THE INVENTION

This application is a continuation-in-part of applicant's copending Ser.No. 07/891,167, filed May 29, 1992, now U.S. Pat. No. 5,273,725, and theinvention concerns the processing of various mineral-containingmaterials, i.e., mineral ores or the processing tailings thereof such asrelatively insoluble residues or the solubilized raffinate or wastestreams or the like generated during the production of metals ornon-metals from their ores. The invention particularly concerns theprocessing of such materials in the form of matrices which are difficultto solubilize and contain relatively substantial radionuclide values andfluorine values essentially entrapped therein. The present processing,in one primary embodiment, is intended to diminish markedly theradioactivity levels in useful non-radionuclide product or in finaldiscard derived from the present processing, which product can be safelyhandled by subsequent users, and which discard can meet chemicallandfill quality standards, i.e., non-hazardous status. The presentinvention is especially adapted to isolate fluorine values, e.g., asfluoride values, and radionuclide values for recovery, and isparticularly suited to the efficient treatment of calcium fluoridesediments derived from ore treatment tailings or wastes which havereceived one or more HF digestions or other treatments followed by limeneutralization.

BACKGROUND OF THE INVENTION

The production of many metals or useful metal-containing compounds fromores, for example, tin from cassiterite, tantalum and niobium fromtantalite/columbite ores, scandium from thortveitate, uranium frompitchblend or uraninite, and rare-earth elements from monazite orbastnasite, typically employs at some point in the production process orin the treatment of tailings therefrom, a hydrofluoric acid digestion,leach or the like. Such treatment is often necessary to convert themyriad of different components including insoluble minerals, metals orcompounds thereof, including various radionuclides, such as theiroxides, halides, carbonates, fluorides, phosphates, sulfates or the liketo species which are soluble in aqueous systems such that they can beseparated out by selective precipitations, extractions or the like. Theinsolubilities of these components are often magnified by the refractoryor inaccessible crystalline nature or the material, i.e., the matrix inwhich these components are occluded, entrapped or chemicallyincorporated, thus necessitating the drastic and extensive HF treatment.Also, in many such processes, HF or other fluorine containing reactantis employed for converting one species to another, such as theconversion of various oxides of uranium to the highly soluble UF₆. Theresult of these HF treatments is that somewhere in the processingoperations, whether primary ore manufacture or tailings processing, themassive and extremely deleterious HF levels must be neutralized,typically and conveniently with lime. Such neutralization, of course,produces large quantities of insoluble calcium fluoride waste sedimentor sludge which either physically entraps or chemically incorporatesinto its crystal lattice residual but significant quantities of valuablemetals including radionuclides such as uranium and thorium, as well asbringing down large and valuable fluorine values. The typical fluorinevalues in these sediments by weight, dry basis, are from about 35% toabout 45%, and for each of uranium, thorium, tantalum, and niobium about0.1%. Greater or lesser concentrations of these elements, or differentmetal species may be found depending on the type of metal ore andprocessing operation involved. It is noted that the quantities of HFneeded in various ore treatment operations around the world is enormousas well as expensive, and thus, fluorine values recovery is a veryimportant consideration in any process directed to the treatment of ore,per se, or of ore processing tailings or waste material or streams.

In this regard many ores, per se, depending on their geographicalsources, as well as the residual wastes from their processing intometals or compounds thereof naturally contain substantial fluorinevalues, for example, fluorite, cryolite, fluorapatite, sellalite, topaz,villiaumite and bastnasite contain fluorine values ranging up to about50% by weight, and often contain substantial radionuclide values. It isnoted that in many instances the initial ore may contain only relativelylow concentrations of radionuclide materials but during processing ofthe ore by flotation, sedimentation, extraction, chemical precipitation,evaporation or the like operations, the level of radionuclide materialsand also of fluorine values can become concentrated and raisedsubstantially, e.g., assaying at least above about 30 pCi/g, and oftenmuch higher.

In recovering the various metal and non-metal components therefrom, theore is typically digested in an acid such as H₂ SO₄, HCl, HNO₃ or evenHF and substantial fluorine values are liberated, usually as HF. Thus,the recovery of these fluorine values becomes an important considerationwhether the values are naturally occurring in the ore or are man-made;but, incident thereto, a two-fold problem, discussed in detail below,exist. These problems which are associated with ore processing wastetreatment operations derive from the fact that the fluorine values mustbe recovered in an economical manner while concurrently reducing theradionuclide levels therein and in final residue or discard to chemicallandfill quality, and also, preferably while recovering the radionuclidevalues, particularly uranium.

For the ore treatment scenario discussed above, certain valuable metalores come into the production plant with substantial radionuclidevalues, for example, tantalum and niobium-bearing materials may containup to about 1.0% by weight uranium and thorium, with associated daughterproducts in equilibrium. When the ores are digested, for example, withHF, the various radionuclide values typically do not mobilize into theliquor in an efficient and consistent manner. Therefore, the final oreprocessing residues or tailings contain substantial quantities ofuranium and thorium as insoluble fluorides, and other insoluble metalfluorides such as scandium. The relatively high radionuclide contentcreates a large problem to the disposal of such residues in that theirclassification according to Federal Regulations is not chemical wastebut hazardous nuclear waste. Typically, the fluoride content in theaforementioned ore residues can range up to about 30% by weight orhigher. Other typical metals which are present in the residues include:calcium, 15%, aluminum, 5%; zirconium, 2%; tantalum, 1.2%; niobium,1.7%; and scandium, 0.15%, the percentages being on dry basis. Thesefluoride-bearing ore residues which often having radioactivity levels ofwell above 30 pCi/g, are usually impounded at the plant site until adetermination can be made concerning their disposition.

As mentioned above and as is discussed hereinafter in greater detail,the relatively insoluble mineral fluoride matrices resulting from thetreatment of ore residues and sediments described above typically havebeen treated heretofore with sulfuric acid at elevated temperatures. Inthis process, hydrofluoric acid is formed and liberated. With thefluoride removed as HF, the various metals within the matrix areconverted to soluble sulfates which can then be dissolved in an aqueoussolution and removed by extraction, precipitation, or other techniques.The purpose of such metal recovery is to generate revenue, e.g., fromscandium, yttrium, or other valuable metals or to decontaminate solids,i.e., remove uranium, thorium or radium to obviate the premium chargedfor radioactive waste disposal.

DISCUSSION OF THE PRIOR ART

Heretofore, an integrated system dedicated to solving the aforestatedtwo-fold problem has not been directly addressed as will become evidentfrom the following discussion concerning specific prior processingsystems for ore or for ore waste or tailings.

In U.S. Pat. No. 5,023,059 a process is disclosed for the recovery ofmetal values such as columbium (Nb), tantalum, thorium and uranium. Therecovery of the Ta, Nb metal values is the first step of the process anda digestion of the sludge is made with HF to isolate the Ta, Nb valuesin a liquid stream from insoluble digest residues which contain the U,Th and other metal values. Consequently, the digest residue, asmentioned above, is roasted with H₂ SO₄ to solubilize the U and Thvalues. This roasting also releases gaseous HF which must be handledquickly, with great care and at great expense. Also, experience hasshown that for refractory CaF₂ type matrices, the H₂ SO₄ treatmenttypically leaves substantial uranium and/or other radionuclide values inthe digest residue. It becomes apparent therefore, that the principalobjectives of this prior procedure are quite different from applicant'sand, as will become more evident hereinafter, necessitate the use ofprocedures and steps such as the onerous HF recovery which applicant'snovel process has obviated.

In U.S. Pat. No. 5,084,253 the metal niobium is separated from uraniumby means of a mineral acid digestion of an alloy of the metals, whereinfluoroboric acid is added to facilitate U⁺⁴ production using lowerconcentrations of HCl. After the insoluble niobium oxides have beenseparated from the digest liquor, the U⁺⁴ ion is precipitated with HF togive the insoluble UF₄ green cake. This process does not have to beconcerned with the problems associated with a substantially concurrentrecovery of uranium and fluorine values from a common waste material orore source, and thus, can accommodate the green cake formation. It isparticularly noted that the process of this patent takes measures toavoid the formation of the uranyl ion, the presence of which as will beseen hereinafter, is important to Applicant's process.

Principal objects, therefore, of the present invention are to provide amarkedly simplified process for the recovery of substantial fluorine andradionuclide values, particularly uranium and thorium values, from oreleach residues or the like, wherein the need for handling gaseous HF isobviated and wherein substantially all of said values are substantiallyconcurrently released to the digest liquor; to provide such processwhich is exceptionally efficient in removing radionuclide values fromore processing waste material such that chemical waste rather thannuclear waste status is achieved for the final discard; to provide suchprocess which readily provides a feed material of high fluorine contentand very low radionuclide content for conversion of the fluorine tonon-gaseous, purified chemical intermediate materials; and to provide animproved process for the production of AlF₃ from fluorine containingores or their processing tailings or wastes.

SUMMARY OF THE INVENTION

These and other objects hereinafter appearing have been attained inaccordance with the present invention which, in a principal embodiment,is defined as a process for converting feed materials of high mineralcontent and substantial radioactivity levels to concentratedradionuclide products of high radioactivity levels and to other productsor discard of very low radioactivity levels, wherein said feed materialscontain substantial fluorine values and radionuclide values assayingabove about 30 pCi/g, said process comprising contacting said feedmaterials with an acidic aqueous digest medium containing boric acid andan acid component comprising one or more other mineral acids, the ratioof said medium (L) to said feed materials (Kg) or dry basis ranging fromabout 1/1 to about 20/1, preferably from about 2/1 to about 6/1, thetotal concentrations of all of said acids being sufficient to solubilizesaid values at commercially acceptable rates, the molar ratio of saidacid component to said boric acid charged to said digest medium beingfrom about 20 to about 0.5, said contacting being carried out for aperiod of time sufficient to solubilize said values into a digestliquor, separating said radionuclide values from said digest liquor togive concentrated, substantially purified radionuclide values, andcontacting the resulting first raffinate liquor with cation reactant toform and precipitate fluoride salt product, said salt product, theresulting second raffinate liquor, and said first raffinate liquor eachhaving a radionuclide values assay of less than about 20% of theradioactivity level of said feed material.

In certain preferred embodiments:

(1) said feed materials comprise ore leach residues containingsubstantial calcium fluoride matrices, the said acid component containssulfuric acid in sufficient quantities to precipitate substantially allcalcium ions as digest residue, wherein said digest liquor is separatedfrom said residue, and wherein said residue has a radionuclide valuesassay of less than about 30 p Ci/g;

(2) said feed materials contain substantial metal fluoride values,wherein said acid component comprises one or both of nitric acid andhydrochloric acid and wherein said first raffinate liquor is contactedwith sulfuric acid to form and precipitate metal sulfates having aradionuclide values assay of less than about 30 p Ci/g;

(3) the metals of said sulfates of (2) above comprise at least calcium,zirconium, iron and manganese.

(4) the molarity of each of said boric acid and said acid component isabove about 0.5, and the aqueous system is heated to a temperature offrom about 75° C. to about 100° C. for a period of from about 0.5 toabout 10.0 hours;

(5) the molarity of said acid component is higher than the molarity ofsaid boric acid;

(6) the said feed materials are redigested one or more times in freshacidic aqueous systems to solubilize the more intractable values;

(7) the initial molar ratio of other mineral acid to boric acid chargedto said acidic aqueous system is from about 1.0 to about 10;

(8) the solubilized uranium values are separated from said digest liquorby passing said liquor through a chelating ion exchange column;

(9) the process of above embodiment (1) wherein said feed materialscontain significant tantalum and/or niobium values, the said digestmedium contains sulfuric acid/boric acid in sufficient quantities toproduce a residue concentrated in tantalum and niobium values, andwherein the residue contains reduced radionuclide values, most notablyuranium and thorium.

BRIEF DESCRIPTION OF THE DRAWINGS

The invention will be further understood from the drawings herein,wherein:

FIG. 1 is a schematic flowsheet for the present overall process;

FIG. 2 is a schematic flowsheet of preferred embodiments of fluorinevalues recovery; and

FIG. 3 is a graph which compares the uranium values leachingeffectiveness of various digest agents versus the present mineralacid/boric acid, digest medium.

DETAILED DESCRIPTION OF THE INVENTION

As indicated above, the invention concerns a process especially suitedfor removing valuable metal, fluoride, and radionuclide values from afeed material of high mineral content in which the metals andradionuclides are present as substantially water insoluble fluorides orare trapped within a metal fluorine matrix which is substantiallyinsoluble in chemical reactant systems. FIG. 1 is a basic, greatlysimplified, but entirely operable schematic for the present overallprocess in one of its preferred embodiments. The process as shown inFIG. 1 comprises a mineral acid/boric acid leach or digestion wherein atleast the fluorides and radionuclides are placed in solution in a digestmedium, a filtration system for removing any remaining insolublematerial, a separation step where solubilized metal values, radionuclidevalues, and contaminants are removed by specific extraction,precipitation, ion-exchange, or the like procedures, and a basic digestof the first raffinate liquor wherein the fluorine values are convertedto fluoride product precipitate, a filtration system for recovery ofsaid fluoride product, and if desired recycle of the second raffinateliquor containing recycle H₃ BO₃ back to the digest medium.

The present process can be applied to the efficient recovery ofstrategic metals including, but not limited to, transition, lanthanide,and actinide series metals from metal-fluoride matrices. Examples ofthese matrices include calcium fluoride sediments from wastewatertreatment operations or from leach ore residues which contain calciumaluminum fluoride materials from primary metal production operations.The focus of the present invention is on the leaching or dissolutionreaction in which the feed materials are reacted with chemical reagentsunder controlled conditions of time and temperature. These chemicalreagents are selected to enhance the effective dissolution of the matrixdirected to optimizing the release of certain metal species andproviding a resulting liquor that may be treated for the efficientrecovery of metal values and fluoride or fluorine values.

Typically in the past, metal-bearing fluoride compounds resulting fromresidues and sediments described above have been treated with mineralacids such as H₂ SO₄, at high temperatures to dissolve or mobilizemetals existing as insoluble metal-fluorides. Recovery of these metalsis desirable for revenue value or for the purpose of decontaminatingsolids, i.e., the removal of uranium and thorium, in order to eliminatethe premium charged for radioactive waste disposal. For example, whencalcium fluoride sediments are reacted with sulfuric acid, a calciumsulfate (gypsum) material and a metal-rich HF/H₂ SO₄ liquor areproduced. The liquor is first separated from the solids and thenprocessed to extract the metal species. The resulting residual solidsare evaluated for chemical content and either redigested to removeaddition metal species or discarded. Typically, such a digestion leavesconsiderable radionuclide values in the residual solids.

The present process uses boric acid as a strong complexing agent in asulfuric acid leaching system for the digestion of a metal fluoridematrix. Preferably stoichiometric amounts of sulfuric and an excess ofboric acid contact the metal fluoride matrix within fairly wide rangesof temperature and reaction times. The boric acid apparently forms avery stable complex with the fluorine in the matrix during the digest orleaching reaction, for example, with calcium fluoride according to thereaction

(1) H₂ O+2CaF₂ +H₃ BO₃ +2H₂ SO₄ →2CaSO₄.2H₂ O+HBF₄.

FIG. 3 hereof provides a comparison of leaching efficiencies for variouschemical reagents alone versus the present complexation with boric acid,particularly, H₂ SO₄ /H₃ BO₃.

After the desired metals are extracted from the resulting HBF₄containing liquor, particularly the radionuclides, e.g. usingion-exchange, solvent extraction, or selective precipitation, theresultant raffinate liquor may then be treated to recover fluorinevalues. This can be achieved, for example, by the preferred end reaction

(2) 12H₂ O+3(NH₄)BF₄ +4Al(OH)₃ →4(AlF₃.3H₂ O)+H₃ BO₃ +3NH₄ OH, or by theend reaction

(3) 3HBF₄ +Al₂ O₃ 3H₂ O+3Na₂ CO₃ →2AlF₃.3NaF+3H₃ BO₃ +3CO₂ which formscryolite, a major feed material used in the manufacture of aluminum.Other aluminum fluoride salts, or the AlF₃ may be formed by similarreactions, e.g., by the reaction with Al₂ (SO₄)₃.

With respect to the digest process, several factors influence the choiceof the mineral acid, including the chemical composition of materials tobe leached, the type of separation system to be used to remove metalvalues or contaminants from the leach liquor, and values recovery cost.In this regard, the mineral acid must form soluble salts with thecomponents of interest and be compatible with the extraction or otherseparation system of choice. For example, sulfuric acid is typicallyused if uranium is the component of interest because it forms a solublesalt, is low in cost, and is compatible with several extractiontechniques. If radium is the component of interest, radium formsinsoluble sulfate salts, and therefore, hydrochloric or nitric acidwould be the mineral acid of choice.

The amounts of mineral and boric acids will vary with the feedcomposition. The mineral acid is typically added in sufficient excess totransmute to soluble species any hydroxides or carbonates present in themineral feed material, to achieve an acid molarity of at least about onenormal, and to neutralize hydroxyl ions formed from the complexationreaction of the fluoride species with the boric acid. The complexationreaction proceeds according to the equation, H₂ O+2CaF₂ +H₃ BO₃ +2H₂ SO₄→HBF₄ +2CaSO₄.2H₂ O. A 25-50% excess of the stoichiometric amounts ofboric acid required to complex the fluoride in the mineral materialpreferably is used and sufficient water is added to maintain the boricacid concentration in the one molar range. A pH<1 is preferred for thisreaction. This preferred molarity is based on the solubility of boricacid at room temperature. Less water can be utilized if sufficienttemperatures are maintained throughout the process to maintain the boricacid in solution.

The mineral materials feed is typically leached for two to four hours atabout 80°-95° C., although temperatures of from about 60° C. or lower,to about 115° C. or higher are useful. Dependent on the feed materialcomposition, leaching times of up to twelve or more hours may berequired, and for particularly refractory materials, multiple leachesmay be expedient.

Following the leaching or digestion process, the leach liquor isfiltered to remove any insoluble components. The solids are washed withwater or with acidified water to remove any residual mineral acid andboric acid from the filter cake. The wash is combined with the filtrateand the liquor is sent to one or more separation operations for theselective removal of metal values or contaminants. The types ofseparation operations employed will vary depending on the mineral feedmaterial, the mineral acid used, and the metals or radionuclides to berecovered. However, in all cases, an ammonium hydroxide precipitation ofcontaminant metals is preferred and is performed prior to the aluminumreactant digestion used for production of aluminum fluoride, as shownabove.

Following removal of metal contaminants, the process liquor, in oneembodiment of AlF₃ production as shown in FIG. 2, is digested for 30minutes to 1 hour with aluminum hydroxide under pressure, at about 150°C. The aluminum hydroxide is added as a solid at about 95% of thestoichiometric quantity required per the equation, 12H₂ O+4Al(OH)₃ +3NH₄BF₄ →4AlF₃.3H₂ O+3H₃ BO₃ +3NH₄ OH. After digestion, the aluminumfluoride product is filtered, the filter cake is washed with water toremove residual ammonium salts and boric acid, the wash is combined withthe filtrate, the resultant liquor is returned to the leaching process,and the aluminum fluoride filter cake is calcined at about 250° C. toremove the waters of hydration and provide a more acceptable product forsale.

The above FIGS. 1 and 2 show the principal steps of digestion,radionuclide separation, and fluoride values recovery and their sequencein the present processes, and also indicate at what stage in the processthe first, second, third, etc., raffinate liquors referred to in theclaims occur. In this regard, the numerical designations of theseraffinate liquors do not signify that they are the only raffinates thatcan and will be produced in carrying out the present process on acommercial basis and with commercial equipment. These designations do,however, represent principal stages which constitutes the presentprocess. For example, formation of the first raffinate liquor, i.e., thesolution remaining after separation of the radionuclides from the digestliquor can be preceeded by any number and composition of secondaryraffinate liquors which may result from any number and types ofsecondary treatments of the digest liquor prior to the principal step ofseparating out said radionuclide values. Such secondary treatments,e.g., may comprise specific chemical precipitations, solventextractions, aqueous or chemical washings, filtrations, centrifugations,settlings, or the like designed to remove specific elements compounds orimpurities from the digest liquor. A specific example is the optionalsecondary removal of zirconium from the digest liquor by means ofammonium hydroxide precipitation prior to separation of the radionuclidematerials from the liquor. Such a precipitation of zirconium compoundwould produce a secondary raffinate, however, the principal character ofthe digest liquor, i.e., as containing the principal fluorine andradionuclide values would not be altered and the designation of "firstraffinate liquor" would still be viable since it resulted from digestliquor containing the said principal values of interest. This sameanalysis or characterization holds also for the other designatedprincipal raffinates and the aforementioned secondary treatments may beemployed, if desired, anywhere along the process, e.g., as preparatoryto or subsequent to precipitation of fluoride salt product.

The various ores, per se, to which the present process is applicableinclude, for example, a typical tantalite ore which may contain fromabout 20% to about 50% tantalum and from about 1% to about 10%columbium, and a columbite ore which may contain from about 3% to about20% tantalum and from about 10% to about 30% columbium. Eastern(Malaysian) tin slags may contain, e.g., 2% to 5% of each metal, whileAfrican tin slags may contain, e.g., 5% to 15% each. Impurities in tinslags typically include CaO (5-20%), SiO₂ (10-40%), FeO (5-20%), TiO₂(5-15%), Al₂ O₃ (5-10%), ZrO₂ (1-5%), and various other minor elementsand usually include significant thorium and uranium values.

A specific example of a feed material useful for the present inventionare actual tantalum ore tailings, the residue analyses of which, asdetermined by neutron activation analysis (NAA), inductively coupledplasmaspectrophotometry (ICP), and wet chemistry techniques are shown inTable I. An X-ray diffraction analysis (XRD) of the tailings showed theprominent compounds to be Ca₁₂ Al₂ Si₄ (SO₄)₃ F₄₀ and Ca₃ Al₂ (Ce)SO₄F₁₃, with CaF₂ and SiO₂ also being detectable. Other elements in traceamounts typically are present also.

                  TABLE 1                                                         ______________________________________                                               Element                                                                              Percent                                                         ______________________________________                                               Al     4.7                                                                    Ca     14.7                                                                   Ce     0.95                                                                   F      32.7                                                                   Fe     1.02                                                                   Mg     1.05                                                                   Mn     0.35                                                                   Na     1.96                                                                   Nb     1.68                                                                   Ni     0.06                                                                   Pb     0.05                                                                   Sc     0.24                                                                   Si     2.0                                                                    SO.sub.4                                                                             3.6                                                                    Ta     1.59                                                                   Ti     1.99                                                                   Th     0.50                                                                   U      0.12                                                                   V      0.11                                                                   Zn     0.02                                                                   Zr     1.78                                                            ______________________________________                                    

The sludges remaining after HF digestion of the ores, slags orconcentrates are primarily unreacted materials, insoluble fluoridesalts, and hydrofluoric acid-containing digestion liquors. A typicalsludge stream remaining after digestion contains about 30% to 60%moisture and, on a dry-weight basis, up to about 3.6% tantalum, up toabout 3.0% columbium, up to about 1% uranium, up to about 1% thorium,and up to about 40% fluorides. On dry basis, the sludge weight is about25% to 75% of the initial input solids weight depending on the origin ofthe feed materials.

The following examples will further illustrate practice of certainembodiments of the invention, wherein all percentages are by weightunless otherwise specified and all metal concentrations are based onelemental weight thereof.

EXAMPLE A Processing of Uranium-Contaminated Calcium Fluoride

A calcium fluoride sludge resulting from the treatment of the processeffluent stream from a UF₆ production facility with lime was processedin accordance with the present invention, the stated componentpercentages or concentrations being approximate. The sludge consisted of40% water and 60% solid material. The composition of the solid materialwas 90-95% calcium fluoride, 5% lime and trace metal hydroxides, and1600 ppm U. The sludge was leached with a 5% excess of sulfuric acid anda 25% excess of boric acid. The purpose of the leach was to extract thefluoride and uranium from the sludge and leave a calcium sulfate leachresidue that could be landfilled.

In carrying out the above, a quantity of 140 g of calcium fluoridesludge was digested in a solution of 105 g of concentrated sulfuricacid, 300 ml of water, and 38 g of boric acid for 3 hours at 90°-95° C.The digest solution was filtered and the calcium sulfate filter cake waswashed with two 50 ml aliquots of water. The filter cake was dried at105° C. and assayed for water and ppm uranium. The cake contained 43%water and 15 ppm (10 pCi/g) U on a dry weight basis. Followingfiltration, the digest liquor was diluted and analyzed. This liquorcontained approximately 139 mg/L of U, 10 g/L of sulfate, and 6.68 g/Lof boron as H₃ BO₃. Based on these results, 99+% of the U was digested,90% of the sulfuric acid was consumed, and 99+% of the boric acid wasrecovered.

EXAMPLE B Benefaction of Ta/Nb Ore Residues

Tantalum and niobium ore residues containing, e.g., 1-2 wt % of eachcomponent may be benefited by the present process. Typically, thetantalum and niobium in the ore residues are contained in a refractiveiron/manganese matrix. This, plus the fact that the metals are at lowconcentrations, makes direct processing economically unattractive. Thepresent strong mineral acid/boric acid combination attacks the insolubleradionuclide fluorides and mobilizes them into the digest liquor wherethey can be extracted, or otherwise isolated, thereby providing aresidue with markedly reduced radionuclide values. In addition, theresidue is beneficiated in tantalum and niobium concentrations in that alarge fraction of other components are efficiently solubilized into thedigest liquor, leaving behind, in the residue, tantalum and niobiumconcentrations that are approximately 10-fold greater than in thestarting material.

While sulfuric acid/boric acid may be used, the selection of nitric acidand boric acid is preferred when calcium and radium need to be removedfrom the ore residues. Nitric acid/boric acid combination is thepreferred selection to provide a residue with lower radioactivity andhigher tantalum and niobium values than normally achievable withsulfuric acid/boric acid combination. In this regard, in a pure acidsystem, some fraction of the tantalum and niobium in thefluoride-bearing residues will be mobilized to the digest liquor asfluoride complexes. The use of sulfuric acid alone mobilizes a largefraction of tantalum and niobium in some sludges, but only a minimalconcentration of tantalum and niobium in others. Since boron reactsstrongly with fluoride in the sludge, the tantalum and niobium cannotreadily mobilize as soluble fluorides and therefore remain as insolublecompounds in the ore residue as a result of the present process.

Processing of Ta/Nb Ore Processing Residue

A sludge (wet residue) resulting from the digestion of a Ta/Nb ore in HFand sulfuric acid was processed in accordance with the present inventionusing a two-stage digest. The sludge contained approximately 52% waterand the dried sludge (residue) consisted approximately of 20-25% F, 15%Ca, 6% Al, 5% Zr, 3% Ti, 2% Fe, 1% Ta, 1% Nb, 0.6% Mn, 0.4% Th, 0.2% U,0.2% Sc, and contained 400 pCi/g Ra-226. The Ta and Nb were present inthe sludge in a refractory iron/manganese containing matrix. The bulk ofthe remaining sludge components were present as insoluble fluoridespecies.

Because of the insolubility of the calcium and radium sulfate, nitricacid was selected as most appropriate for the aforesaid other mineralacid. In accordance with the present invention, 210 g of the raw sludgewas digested in a two-stage, countercurrent, continuous fashion with 120ml of concentrated nitric acid, 30 g of boric acid, and 300 ml of water.The digestion times and temperatures for each of the two stages were 4hours and 80°-90° C., respectively. In typical countercurrent fashion,the digest residues left from the first stage were contacted with freshdigest solution in a second stage. The second digest solution was thenused to contact the raw sludge. Residues from the first digest weretransferred to the second digest vessel by filtering the digest leachsolution and transferring the filter cake to the second digest vessel.Following the second digest, the second digest solution was filtered andthe digest residues were washed with approximately 50 ml of 0.1 molarnitric acid. The wash solution was combined with the digest solution andboth were transferred to the first digest vessel along with 210 g of rawsludge.

After completion of the second digest, the residues thereof were driedand assayed. These residues typically weighed approximately 5 grams andconsisted of approximately 50% Ta/Nb oxides. Essentially 95% of theremaining sludge constituents were solubilized into the first digestliquor including the Sc, Th, U, and Ra-226. The combined digest and washliquors counted about: 200 pCi/ml total uranium; 25 pCi/ml thorium; 70pCi/ml radium-226; and contained 230 mg/L of scandium.

An encapsulation of preferred technology as it relates specifically totantalum and/or niobium recovery from residues which containeconomically attractive amounts of either of these metals is as follows:

The process for recovering Ta and/or Nb values from highly fluorinatedore materials comprising;

(a) contacting said materials with an acid component comprising HNO₃ orHCl or mixtures thereof, and H₃ BO₃ wherein the H₃ BO₃ to acid componentmolar ratio is from about 1/10 to about 1/1,

(b) maintaining the temperature between about 55 C. and about 85 C.,

(c) continuing the reaction for a sufficient period to substantiallysolubilize the major portion of said materials and to substantiallyinsolubilize all Ta and/or Nb values, and

(d) separating the solids containing the Ta and/or Nb values fromsolution.

Preferably in the process the separated solids from step (d) aresubjected to a reducing environment such as hydrogen. Also preferred isthe use of the nitric acid-boric acid combination.

EXAMPLE C Uranium Extraction and Recovery from Calcium Fluoride DigestLiquors

The liquor and combined washes obtained from the digestion of calciumfluoride sludge illustrated in Example A were used in this experimentand contained approximately 130 mg/L total uranium as a solutionapproximately 0.4N in fluoroborate ion at a pH of about 0.5. It is notedthat in carrying out the present process, the initial digest pH istypically between about 0.0 and about 3.0, and preferably is from about0.5 to about 1.5. Extraction of the uranium was performed by using animinodiacetate ion-exchange resin SR-5, manufactured by Sybron/IonacChemicals. Ten cc of SR-5 with standard particle size distribution-16+50 Standard U.S. Mesh were loaded into a laboratory column. Thiscolumn was treated with 1M sulfuric acid and subsequently rinsed untilthe effluent had reached a pH of 2.6. The uranium-bearing digestsolutions were adjusted to pH of 2.8 with 10N sodium hydroxide and fedto the column at an average flow-rate of four bed volumes (BV) per hour.The column effluent was periodically sampled and analyzed for totaluranium by a laser induced phosphorescence method as described in anarticle by B. A. Bushaw, entitled "Progress In Analytical Spectroscopy",proceedings vol. 26, Oak Ridge Conference On Analytical Chemistry InEngineering Technology, Knoxville, Tenn., October 1983. Through fourteenBV of solution processed, the leakage of uranium in the effluent was<3%, and through 30 BV of liquor processed, approximately 91% of uraniumin the feed solution, i.e., 36.0 mg of radionuclide, was extracted bythe SR-5.

Uranium was then recovered from the loaded laboratory column. For thisrecovery, the column was first rinsed with 20 ml of deionized water todisplace uranium from the void volume. The elution of uranium from theion-exchange resin was performed by passing a 2N nitric acid solutionthrough the column and collecting the eluant in 20 ml increments atabout 4 BV/hr. The total uranium analysis in the eluant fraction wasperformed by a laser induced phosphorescence method. Analysis of a totalof eight BV of eluant collected showed that essentially 100% of theuranium values were stripped from the column.

Methods For Radiometric Analyses

For the analyses of radionuclides, two types of methods exist i.e.,destructive and non-destructive techniques. For the destructivetechnique, the sample matrix is converted to a form where the targetnuclide may be selectively isolated and analyzed. For thenon-destructive technique, the sample matrix is assayed for targetnuclides without destroying or converting the matrix.

Uranium was assayed in several ways. For the calcium fluoride testing,when uranium was in a solid matrix, this matrix was digested and theresulting digest solution was analyzed by kinetic phosphorescence. Inthis technique, a chemical is added to the liquid to complex with theuranium. This complex is excited with a laser which causes it tophosphoresce. The decay of the intensity of the phosphorescing speciesor complex is followed and plotted over time. Extrapolation of the curveto t=0 provides the concentration of the uranium. When liquids areanalyzed, as from ion-exchange processing or in the mother liquor, thedye is added directly.

Thorium was measured by Inductively Coupled Plasma Spectroscopy (ICP).Thorium measurements in solids, e.g. digest residues, first requirescomplete digestion of the residue to produce a liquid, and the thoriumanalyses in the liquor is performed directly by ICP. Uranium in leachresidues has also been measured by ICP following acid digestion of theresidues.

Radium analyses were performed by gamma spectroscopy. In thisnon-destructive method, the gamma rays emitted from either the radium orfrom its daughter products are collected and assayed and then related tothe total radium concentration. Pre-treatment of the sample prior toassay is not required by this measurement technique.

Other useful analytical methods include:

Metals--Atomic Adsorption or ICP;

Fluorides--Selective Ion Electrode;

Fluoroborates--Selective Ion Electrode or Gravimetric using limeaddition;

Nitrates--Selection Ion Electrode.

Ionac SR-5 is a macroporous chelating resin based on iminodiacetic acidfunctionally for the selective recovery of heavy metal ions from wastetreatment streams and process liquids. The resin is designed to stand upto the harsh conditions of waste and process applications. Resinattrition is minimal due to the rugged, second-generation macroporouspolymer backbone and the very low (approximately 20%) reversibleswelling. The general specifications of this resin are as follows:

    ______________________________________                                        Polymer Structure   Styrene-Divinylbenzene                                                        Copolymer                                                 Functional Groups   Iminodiacetic Acid                                        Physical Form       Moist Spherical Beads                                     Ionic Form (as shipped)                                                                           Sodium                                                    Selective Copper Capacity (Na form)                                                               1.7 moles/l                                               Reversible Swelling (H to Na)                                                                     20% (approx.)                                             Shipping Weight     45 lbs/cu. ft (720 g/l)                                   Particle Size Distribution (wet)                                               (U.S. Std. Mesh)   -16 + 50                                                   (Metric)           0.3-1.2 mm                                                Effective Size (mm) 0.47-0.53                                                 Uniformity Coefficient                                                                            1.7 Maximum                                               Water Retention     45-50%                                                    pH Range (Stability)                                                                              1-14                                                      Solubility          Insoluble in all common                                                       solvents                                                  ______________________________________                                    

Further details of this resin are found in the sales literature, 5pages, of Sybron Chemicals Inc., of Birmingham, N.J., entitled IONACSelective Metal Recovery.

EXAMPLE D Recovery of Fluorine Values As Aluminum Fluoride

Three processes for the preparation of high-quality aluminum fluoridefrom calcium fluoride leach liquors as prepared above, were developed.These processes follow the general flowsheet presented in FIG. 2. It isparticularly noted that heretofore, aluminum trifluoride has beenproduced mainly by the reaction of HF with aluminum hydroxide, whichprocess requires the very costly HF reactant as well as theextraordinary anti-corrosion equipment and handling precautions requiredfor and associated with HF.

In the present process #1, aluminum is added as alumina trihydrate,Al(OH)₃, to leach liquor from which incidental or trace metallicimpurities have been removed by precipitation with ammonium hydroxide.The Al(OH)₃ digestion is carried out at a temperature of from about 150°C. to about 195° C., preferably about 180° C. and a pressure of fromabout 180 to about 215 psi, preferably at about 200 psi for about 20-60minutes, preferably about thirty minutes in a teflon bomb.

In the present process #2, after removal of incidental or trace metallicimpurities from the leach liquor by hydroxide precipitation, aluminumsulfate is added to the leach (raffinate) liquor and the digestioncarried out for about 2-8 hours, preferably for approximately 4 hours atabout 90°-95° C. Ammonium hydroxide is added periodically throughout thedigestion period to neutralize any free acid such as H₂ SO₄ formedduring the reaction of ammonium fluoroborate with aluminum sulfate.

In the present process #3, after removal of incidental or trace metallicimpurities from the leach liquor by hydroxide precipitation, or otherseparation processes such as solvent extraction or ion exchange, asufficient quantity of sulfuric acid is added to adjust the pH of theliquor to between about 1 and about 4, preferably in the 2 to 3 range.Aluminum hydroxide, preferably in stoichiometric amounts, but may be in,e.g., about 20% excess, as alumina trihydrate is added to this liquorand a digestion is carried out from about 3 to about 6 hours at fromabout 80° C. to about 97° C., preferably from about 90° C. to about 95°C. The resulting slurry is filtered to provide an aluminum fluoridetrihydrate.

The following equations show the end chemical reactions for each processwherein, e.g., up to about 20% excess reactant may be employed, butwherein substantially stoichiometric quantities of the reactants arepreferably employed.

Process #1

12 H₂ O+3(NH₄)BF₄ +4Al(OH)₃ →4AlF₃.3H₂ O+3H₃ BO₃ +3NH₄ OH

Process #2

3(NH₄)BF₄ +2Al₂ (SO₄)₃ +12H₂ O+9NH₄ OH→4AlF₃.3H₂ O+3H₃ BO₃ +6(NH₄)₂ SO₄.

Process #3

9H₂ O+3HBF₄ +4Al(OH)₃ →4AlF₃.3H₂ O+3H₃ BO₃

In carrying out the above processes, the aluminum fluoride is filteredand washed with water to remove boric acid and sulfate contamination.The aluminum fluoride filter cake is then calcined to remove water andis then ground to meet particle-size specifications. The digestionliquor is treated with lime and the ammonium hydroxide recovered. Thecalcium sulfate resulting from this lime treatment is filtered andwashed with water to remove boric acid contamination and the liquor andwash water then returned to the CaF₂ leaching process for reuse. It isnoted that the extremely efficient recovery of the boric acid is a veryvaluable and unexpected aspect of these processes.

Process #1

A synthetic leach liquor was prepared to evaluate process #1. The leachliquor was prepared by adding 91.5 g of 48% fluoroboric acid, 7.73 g ofboric acid, and 25 g of (NH₄)₂ SO₄ to a 500 ml volumetric flask anddiluting to 500 ml with demineralized water. This resulted in asynthetic leach liquor containing 1 molar fluoroboric acid, 50 g/Lammonium sulfate, and 0.25 molar boric acid which is typical of acalcium fluoride leach liquor. A 50 ml aliquot of the synthetic leachliquor was neutralized with approximately 3.3 ml of concentratedammonium hydroxide to pH 7.0 This solution was placed in a 100 ml teflondigestion bomb and 5 g of aluminum hydroxide was added. Only 95% of thestoichiometric quantity of the aluminum hydroxide was added to insurethat aluminum consumption went to completion. The teflon bomb was sealedand the contents were digested by placing the bomb in a microwave andheating for 30 minutes at 150° C., the pressure reaching about 40 psi.The bomb was then heated at 180° C. for another 30 minutes, the pressurereaching 130 psi. After cooling, the contents of the bomb were filtered,the filter cake was washed with approximately 50 ml of water and thecake dried at 105° C. After drying, the cake was analyzed to determineits aluminum and fluoride values. The analytical results showed thatabout 92% of the fluoride was recovered as aluminum fluoride trihydrate.

Process #2

An actual calcium fluoride leach liquor was utilized to evaluate theprocess #2. An aliquot of 920 ml leach liquor containing 0.25 mole/Lfluoroboric acid, 0.08 mole/L boric acid and 0.2 mole/L sulfuric acidwas placed in a one L glass beaker and 87 g of Al₂ (SO₄)₃.18H₂ O wasadded and the solution stirred until all of the aluminum sulfatedissolved. The solution was adjusted approximately to a pH=8 withammonium hydroxide and the beaker placed on a hot plate, stirred, andheated for 4 hours at 90°-95° C. The pH was periodically adjusted duringthe 4 hour digestion period with ammonium hydroxide to maintain a pHgreater then 7.0. Following the digestion period, the solution wasfiltered and the filter cake was washed with approximately 50 ml ofdemineralized water to remove boric acid and ammonium sulfatecontamination. After washing, the filter cake was dried at 105° C. andanalyzed to determine its aluminum and fluoride values. The analyticalresults showed that the filter cake consisted of aluminum fluoridetrihydrate.

An actual calcium fluoride leach liquor was utilized to evaluate theprocess #3. An aliquot of 500 ml of leach liquor containing 0.260 mole/Lfluoroboric acid, 0.11 mole/L boric acid, and 0.11 mole/L sulfuric acidwas treated with concentrated ammonium hydroxide to a pH of near 7 andstirred at ambient temperature for about 15 minutes. The resultingslurry, comprised of precipitated metal hydroxide impurities, wasfiltered and the resulting filter cake was washed with 100 ml ofdeionized water. The washwater was added to the filtrate to bring thetotal liquid volume up to 620 ml.

To this combined volume, concentrated sulfuric acid was added to adjustthe solution pH to about 3. Following pH adjustment, a quantity of 15.1grams of alumina trihydrate (manufactured by ALCOA, Point Comfort, Tex.)were added to the liquor plus washwater solution and heated for 6 hoursat 90° C. This solution was filtered and the resulting filter cake waswashed with 100 ml of deionized water. The filter cake was dried atabout 105° C. and analyzed to determine its aluminum and fluoridevalues. The analytical results showed that the filter cake consisted ofaluminum fluoride trihydrate with ≧90% recovery of fluoride values fromsolution.

The invention has been described in detail with particular reference topreferred embodiments thereof, but it will be understood that variationsand modifications will be effected within the spirit and scope of theinvention.

I claim:
 1. The process for recovering Ta and/or Nb values from highlyflourinated ore materials comprising;(a) contacting said materials withan acid component comprising HNO₃ or HCl or mixtures thereof, and H₃ BO₃wherein the H₃ BO₃ to acid component molar ratio is from about 1/10 toabout 1/1, (b) maintaining the temperature between about 55 C. and about85 C., (c) reacting said materials with said acid for a sufficientperiod to substantially solubilize the major portion of said materialand to substantially insolubilize all Ta and/or Nb values, and (d)separating the solids containing the Ta and/or Nb values from solution.2. The process of claim 1 wherein the solids from step (d) are subjectedto a reducing environment.
 3. The process of claim 1 wherein said acidcomponent is nitric acid.
 4. The process for converting feed materialsof high mineral content and substantial radioactivity levels toconcentrated radionuclide products of high radioactivity levels, toother products or discard of very low radioactivity levels, and torecover tantalum and/or niobium containing compositions wherein saidfeed materials contain substantial fluorine values and radionuclidevalues assaying above about 30 pCi/g, said process comprising contactingsaid feed materials with an acidic aqueous digest medium containingboric acid and an acid component comprising nitric and/or hydrochloricacid, the ratio of said digest medium (L) to said feed materials (Kg) ondry basis ranging from about 1/1 to about 20/1, the total concentrationsof all of said acids being sufficient to solubilize said values atcommercially acceptable rates, said contacting being carried out for aperiod of time sufficient to solubilize said values into a digest liquorand to leave substantial concentrations of said tantalum and/or niobiumcontaining compositions, separating said radionuclide values from saiddigest liquor to give concentrated, substantially purified radionuclidevalues, and contacting the resulting first raffinate liquor with cationreactant to form and precipitate fluoride salt product and to provide asecond raffinate liquor, each of said salt product and second raffinateliquor having a radionuclide values assay of less than about 20% of theradioactivity level of said feed materials.
 5. The process of claim 4wherein said acid component is nitric acid.